Phosphorous pentoxide producing methods and phosphate ore feed agglomerates

ABSTRACT

A phosphorous pentoxide producing method includes forming a reducing kiln bed using feed agglomerates containing a lignin sulfonate both inside the agglomerates and coating a surface of individual agglomerates. Freeboard particulates are generated from the agglomerates, an amount of particulates generated being less than would occur in the method with no lignin sulfonate. Kiln off gas is generated and phosphorous pentoxide is collected from the kiln off gas. The kiln discharges a residue containing processed agglomerates, less than 20% of the agglomerates&#39; phosphate input to the kiln remaining in the residue. The percentage of input phosphate that remains in the residue is less than would occur in the method with no lignin sulfonate. The method may include forming green agglomerates. Controlling a drying rate of the green agglomerates may wick some of the lignin sulfonate onto the surface of individual drying agglomerates without adhering the agglomerates together.

TECHNICAL FIELD

The invention pertains to methods for producing phosphorous pentoxideperformed in counter-current rotary kilns and phosphate ore feedagglomerates for use in such kilns.

BACKGROUND

One known method for producing phosphorous pentoxide (P₂O₅, usuallypresent as the dimer P₄O₁₀ in the gas phase) involves processing rawmaterial agglomerates containing phosphate ore, silica, and coke in thebed of a rotary kiln to chemically reduce the phosphate ore and generategaseous phosphorus metal (P₄) and carbon monoxide (CO) off gas to thekiln freeboard where they are burned (oxidized) with air to provide heatfor the process. It may be referred to as the kiln phosphoric acid (KPA)process. The oxidized phosphorus metal is a phosphorus oxide (normally,P₄O₁₀), which can be scrubbed from the kiln off gases with a phosphoricacid (H₃PO₄) solution and water to make a suitable phosphoric acidproduct.

KPA process chemistry is similar to another process known as the furnaceacid process for manufacture of phosphoric acid. In the furnace process,the raw materials are heated and partially melted. An endothermicreduction reaction is carried out in one vessel called the electricfurnace where the heat is supplied by the use of electric resistanceheating in the bed. The phosphorus metal is recovered from the off gasof the furnace with cold water sprays as liquid phosphorus metal whichcan be transported to another vessel called the burner where itgenerates considerable heat while being burned with air. The resultingphosphorus oxide is absorbed in water to make a concentrated, highpurity phosphoric acid.

The electric furnace in the furnace process does not use the heatgenerated from burning phosphorus metal that arises in the burnervessel. Also, the electric furnace does not use heat from burning thecarbon monoxide that it generates. Although widely used in the lastcentury for producing phosphoric acid, the cost of electricity ascompared to the cost of sulfuric acid resulted in shutdown of most ofthe furnace acid plants in favor of another process known as thesulfuric acid process for making phosphoric acid.

If the heat generated in burning the off gasses from the furnace processreduction reaction could be utilized to provide the heat requirements ofthe reduction process, thereby replacing electrical heating, theneconomies might be realized. Converting the furnace process carbonreductant to carbon dioxide might generate sufficient heat, if usedefficiently, to replace all the heat added by electricity in the furnaceprocess. A vision of such potential motivated many researchers over theyears to develop concepts where heat integration could be realized. Thefollowing references describe the various attempts: Levermore (U.S. Pat.No. 2,075,212), Lapple (U.S. Pat. Nos. 3,235,330 and 3,241,917), Saeman(U.S. Pat. No. 3,558,114), Megy (U.S. Pat. Nos. 4,351,809 and4,351,813), Hard (U.S. Pat. No. 4,389,384), and Park (U.S. Pat. No.4,420,466). All of the described processes use a rotary kiln with areducing bed and an oxidizing freeboard and are collectively within therealm of KPA processes.

A kiln within a kiln concept forwarded in the Levermore patent addressedthe heat integration issue in a reasonable conceptual way, but was notpractical because no material of construction for the inner kiln wasavailable. The Levermore process heated up the agglomerated solidscontaining phosphate ore, silica, and carbonaceous material andconducted the endothermic reduction reaction in an inner kiln heldinside an outer kiln. The P₄ and CO off gases from the reductionreaction in the inner kiln passed between the outside wall of the innerkiln and the inner wall of the outer kiln, where air was admitted tooxidize the P₄ and CO, generating sufficient heat to supply therequirements of the inner kiln. The heat then passed through the outsidewall of the inner kiln.

The Lapple and Megy patents recognized that a separate oxidizingfreeboard and reducing bed could be maintained in a kiln without aseparating wall, but failed because of melting problems during thephosphate reduction reaction. Lapple and Megy specify acalcium-to-silica mole ratio in the feed to the kiln greater than 1.0.

The Park patent describes a process requiring ore with less silica thanis cheaply available and a very hot kiln operation that has not appearedattractive enough to encourage commercialization to date.

The process in the Saeman patent involves carrying out the rotary kilnreduction reaction in a molten slurry, contained within the shell of thekiln protected by freezing a layer of solids on the inside of the kilnwall. This process has been abandoned.

The process in the Hard patent showed promise and a continuous pilotplant based on a 33 inch diameter by 30 feet long kiln was operatedunder the direction of the present inventor in the early 1980's. Theresults were published in Leder, et al., A New Process for TechnicalGrade Phosphoric Acid, Ind. Eng. Chem. Process Des. Dev., 1985, 24,888-897, but the process was abandoned as his teachings were notcomplete enough to show how to carry out an economically feasiblecommercial process. The pilot plant yields were low, with a maximumyield of 72% when run in a commercial mode without bedding coke, and 86%when a large amount of bedding coke was used.

Other problems included: 1) throughput rates that were low whichindicated a high capital cost requirement for the process, 2) hightemperature of operation (greater than 1435° C.) to reach high yields,which put the operation close to melting problems and required highersilica admixtures in the kiln feed than desired for commercialoperation, and 3) high maintenance problems. The high temperature andpartial oxidation of carbon from the kiln solids resulted in transfer ofsignificant amounts of fluorine, sodium, potassium, and sulfur to thekiln off gas, producing deposits in the back end of the kiln and off gaslines, contamination of the product acid, and extra costs associated inscrubbing acidic gases from the process. The off gases from the processwere chemically reducing, requiring the added cost of an after burner.Significant generation of dust was dealt with using dust collectiondevices and kiln ringing from the dust also occurred. The combination ofthese problems was such that no one has attempted to commercialize theHard process even though it has now been over twenty years since thepilot plant was operated.

U.S. Pat. Nos. 7,378,070 and 7,910,080 were issued to the presentinventor for a new process called the Improved Hard Process (IHP). Anumber of process benefits allowed the IHP to be run with an increasedlevel of phosphate ore in the feed, at 120° C. lower operatingtemperature, at higher yields due to avoidance of a competing reactionthat was operable in the Hard process, and with a fivefold increase inkiln phosphorus output for a given size kiln. These process benefitsresulted in breakthrough economic projections, justifying a commercialdemonstration plant. The IHP considerably benefited the capital andoperating cost of running a KPA process.

However, as may be appreciated from the previous difficulties describedabove, the KPA process and/or IHP may be improved further.

BRIEF DESCRIPTION OF THE DRAWINGS

Preferred embodiments of the invention are described below withreference to the following accompanying drawings.

FIG. 1 is a block diagram of the IHP.

FIG. 2 is a chart pertaining to IHP Example 1 showing meltingtemperatures with respect to calcium-to-silica mole ratio and showingthe temperatures for greater than 92% yield in 1, 2, and 4 hours withrespect to calcium-to-silica mole ratio.

FIG. 3 is a chart showing volatility for various elements with respectto temperature.

DETAILED DESCRIPTION OF THE PREFERRED EMBODIMENTS

A further benefit, which is the subject of the present document,involves a new method of providing a carbonaceous coating on the kilnfeed ball and a new feed ball so-coated, which have far reachingpositive effects on the Improved Hard Process (IHP). Details of the IHPapplicable to the embodiments herein are discussed in the section belowentitled “Improved Hard Process” and are provided for the reader tounderstand with greater depth the processes in which the embodimentsherein operate.

Although the IHP provided increased yield over the Hard process andother KPA processes, additional improvement was possible. Carbon burnoutrepresents a significant factor in yield losses for the Hard process,particularly in small kilns, given their small diameter, as explainedbelow. Observation indicated that providing a carbonaceous coating onfeed balls that persisted during appropriate periods of bed heat upprovided measurable increases in yield by further guarding againstcarbon burnout.

The inclusion of lignin sulfonate in feed balls during ball formationwas found to produce such a coating. Essentially, suitable amounts oflignin sulfonate blended with the raw materials for the balls wereobserved to pass, in part, to the ball surface during the dryingprocess. Without being limited to a particular theory, a wickingmechanism enabled by the nature of the drying process is believed toallow formation of the coating. As water moves through the ball to thesurface to evaporate during drying with hot air, lignin sulfonatedissolved into the moving water may remain behind on the surface afterthe water evaporates. The result is a ball with a tough, leather-likecarbon surface that could also be considered akin to a modestly hardrubber.

As the coated ball heats in the kiln, volatiles evolve from the ball,blanketing the ball and reducing ingress of P₄O₁₀ to the ball which, ifit arrives at the ball surface, is believed to back react and formcalcium pyrophosphate. The pyrophosphate would then apparently undergoreduction when the ball reaches reduction temperature, reducing yield,as described further below. Also, without volatiles blanketing anuncoated ball as the ball heats, the carbon burns out of a shell at theball surface by oxidation with freeboard gases. Since the carbon is notavailable in the burned out shell, fluorapatite will not be chemicallyreduced there.

In comparison to balls lacking lignin sulfonate, testing showed that theoverall chemical reduction yield for balls containing lignin sulfonateincrementally increased. Therefore, the testing is believed to supportthe described theoretical mechanism even though not confirmed bycomposition analysis of balls from the bed heat up period.

Notably, volatiles from the lignin sulfonate evolve at a comparably lowtemperature, much of them between the kiln entering temperature of 120°C. and the onset of volatiles from green petroleum coke in the feedballs at 650° C. Beneficially, the lignin sulfonate volatiles evolvejust when they are of greatest benefit, during bed heat up before othervolatiles evolve from other materials. The Megy IHP patents describeusing small additions of coal added among the feed balls as “bedding,”which also have low temperature volatiles to achieve this purpose.However, the lignin sulfonate provides additional benefits.

The fixed carbon component of the lignin sulfonate coating may staybehind between 650° C. and the onset of reduction at 1180° C. In thisregard, it forms a sacrificial layer of carbon right where it is ofgreatest benefit to protect the balls from carbon burnout and thereby toincrease yield. A batch kiln used in testing represents a severe case ofthe carbon burnout due to high turbulence therein. In such case, thelignin sulfonate provided a clear benefit. As the process is scaled upin a larger, continuous kiln, an expectation exists that carbon burnoutwill decrease so that the incremental yield increase might not be sostark. Nevertheless, even a small yield increase may benefit processeconomics.

Perhaps the greatest and most surprising benefit is the suppression ofdusting during the maximum dusting time in the heat up period. Measureddust levels were less than other tests using like feed balls withoutlignin sulfonate additions. The dust levels, though small, in acommercial, countercurrent kiln may lead directly to kiln wall deposits.Reduction in dust level also increases the purity of the phosphoric acidproduct and further reduces wall accretions; a very significant benefit.

As tangential benefit, the lignin sulfonate does not detract from a ballstrength specification. For this metric, balls containing ligninsulfonate were generally incrementally superior to balls containing onlybentonite or clay as a binder. Accordingly, lignin sulfonate may replaceat least part of the binder otherwise used. A residual of clay in theballs may be desirable as it supplies strength in balls after the carbonfrom the lignin sulfonate burns out and in the kiln residue from theprocess when used as a co-product aggregate. Since lignin sulfonate is awaste product, it may be acquired at low cost, but still represents anadded cost, which is best to offset with benefits.

In an embodiment, a phosphorous pentoxide producing method includesforming a reducing kiln bed using feed agglomerates below an oxidizingfreeboard of a counter-current rotary kiln, the bed having a length. Theagglomerates contain phosphate ore particles, carbonaceous materialparticles, sufficient silica particles for the agglomerates to exhibit acalcium plus magnesium-to-silica mole ratio of less than 1.0, and alignin sulfonate both inside the agglomerates and coating a surface ofindividual agglomerates. In the IHP discussed below, the relevant ratiois described as calcium-to-silica. Often, calcium content inagglomerates is much more than magnesium, perhaps by a factor of 50 orso. However, in the interest of fine tuning the process and allowingapplicability to high magnesium content ores, the ratio may be (moles ofcalcium+moles of magnesium)/moles of silica of less than 1.0 or perhapslower, as discussed below.

While ratios between 1.0 and 0.5 can be run in the IHP with high purityore and attention to the details of the IHP specifications, compositionsin this range have a small temperature operating range prior to melting(see FIG. 2), with reduced throughput. When ratios below 0.5 are used,the ball becomes more resistant to melting as the calcium monosilicateproduct from the reduction reaction is formed into long glass-likechains with the addition of a second silica atom. Thus, more robustoperation, with higher throughput and greater tolerance for impurities,is possible with ratios below 0.5.

The method includes maintaining a bed temperature at or above 1180° C.along a portion of the bed length. Freeboard particulates are generatedfrom the agglomerates, an amount of particulates generated being lessthan would occur in the method with no lignin sulfonate. Kiln off gas isgenerated and phosphorous pentoxide is collected from the kiln off gas.The kiln discharges a residue containing processed agglomerates, lessthan 20% of the agglomerates' phosphate input to the kiln remaining inthe residue. The percentage of input phosphate that remains in theresidue is less than would occur in the method with no lignin sulfonate.

By way of example, less than 10% of the agglomerates' phosphate input tothe kiln may remain in the residue. The calcium plus magnesium-to-silicamole ratio may be less than 0.5. The method may further include forminggreen feed agglomerates from a homogeneous blend of water, the ligninsulfonate, the phosphate ore particles, the carbonaceous materialparticles, and the silica particles. The green agglomerates may containabout 0.1 to 4 wt % lignin sulfonate on a dry basis, that is, notaccounting for the presence of water. All compositions expressed hereinfor the feed agglomerates or feed balls are on a dry basis, unlessotherwise indicated. Still further, the green agglomerates may containabout 0.4 to 2 wt % lignin sulfonate. The method may include controllinga drying rate of the green agglomerates and, as the water evaporatesfrom the agglomerates, forming the lignin sulfonate coating onindividual dried agglomerates by wicking some of the lignin sulfonateonto the surface of individual drying agglomerates without adhering theagglomerates together.

The carbonaceous material particles that chemically reduce the phosphateore (not the lignin sulfonate) may contain about 8 to 12 wt % volatilesand at least 80% of all particles may exhibit a size less than 200 mesh.The lignin sulfonate coating may have a thickness of about 1 to 50micrometers. Individual agglomerates may contain a total of about 0.1 to4 wt % lignin sulfonate including the coating. Due to the wickingmechanism, the lignin sulfonate inside the agglomerates may exhibit acontinuously decreasing concentration gradient of lignin sulfonate fromthe surface into individual agglomerates.

The method may further include volatilizing at least a portion of thelignin sulfonate coating during bed heat up from about 650 to about1180° C. Also, the method may further include blanketing the bed withvolatiles evolving at least partly from the lignin sulfonate during bedheat up from about 120 to about 650° C. and retaining a sacrificiallayer of fixed carbon from the lignin sulfonate coating. In thecircumstance where the sacrificial layer is retained past bed heat up toabout 650° C., at least a portion of the sacrificial layer may bevolatilized during further bed heat up from about 650 to about 1180° C.while volatiles from the carbonaceous material particles are alsovolatilizing.

In the IHP, the raw materials are co-ground in an open circuit ball millto provide the 80%-200 mesh according to the methods described below.The open circuit ball mill grind gives a large proportion of thematerial in the 1 to 50 micrometer size range that would not be presentif closed circuit grinding was used. The fineness of the carbon and theintimate mixture both contribute to the IHP kiln reaction going to highyield at 1180° C. With this level of super-fines in the mixture, U.S.Pat. Nos. 7,378,070 and 7,910,080 (the Megy IHP patents) teach that clayadditions help to make sufficiently strong balls. Examples are about0.7% of sodium modified bentonite, or 3% of phosphate clay normallyassociated with phosphate mining operations.

However, as described, use of lignin sulfonate in producing balls alsoshows that lignin sulfonate works well as a binder with about 1.0 wt %addition (dry basis). Unless otherwise indicated, reference herein to acontent of lignin sulfonate is on a dry basis. Lignin sulfonate isavailable in a variety of forms, solid or liquid, including as an about50 wt % viscous, aqueous solution obtainable as a waste product from themanufacture of paper by the sulfite industry. Accordingly, addition ofabout 2.0 wt % of a 50 wt % solution may provide the desired about 1.0wt % of lignin sulfonate. Given that the liquid solution is anindustrial product, actual content of lignin sulfonate may vary from lotto lot and the amount added to provide the desired content may vary. Oneexample of such a product includes calcium lignin sulfonate availablefrom ArrMaz Custom Chemicals in Mulberry, Fla. Several types of ligninsulfonate may be suitable, including ammonium lignin sulfonate, sodiumlignin sulfonate, and calcium lignin sulfonate.

Drying the balls so as to create the mentioned wicking mechanism mayresult in the lignin sulfonate being deposited as an adherent tough skinon the outside of the feed balls. An example of a method of preparingthe feed balls for the kiln is as follows.

The raw materials and a small amount of clay, as discussed above, areproportioned with weigh belt feeders into a drier. After drying the rawmaterials they are co-ground in an open circuit ball mill. Afterco-grinding, the raw materials are mixed with about 8 wt % water and 1.0wt % lignin sulfonate is blended into the dry co-ground raw materials.The moistened, co-ground raw materials are then formed into balls,either on a balling pan or, beneficially, in a balling drum usingprocedures known in the iron ore pelletizing industry. For the ballingdrum, this involves a large recycle stream of the growing balls followedby a roller screen to select those balls that are on size. Althoughballs are described here, the embodiments may be applicable to othergeometric shapes, if desired to meet a specification in a certain IHPprocess.

After suitably sized green balls are made, they are passed to a lowtemperature grate for drying, operating at a temperature of about 250°C. This can be accomplished by passing heated gases either up draft ordown draft (or a combination) through a bed of balls, about 12 inchesthick, on a moving porous belt, which generally dries the balls in about10 minutes. It is during this step that the surprising action of thelignin sulfonate occurs. Decreasing the bed depth to about 6 inches mayallow drying in less time. However, observation indicates that ballintegrity may deteriorate if drying is too fast, for example, 5 minutes.Deterioration may be influenced by heat sensitivity of one or morecomponents. The strength of the composition used for the green ballsdescribed herein allows the increased bed depth, and slower drying,without a concern that the stacking will damage the balls.

Drying temperature may be below about 400° C., at which temperaturecarbonaceous material in the balls may begin to oxidize (burn), or evenless than about 350° C. to be safe. Also, the amount of clay, ifincluded, may influence heat sensitivity. It is believed that too muchclay closes pores in green balls such that steam fissures develop duringrapid drying, weakening the balls. Balls may even explode as a result,if dried too fast. However, clay increases ball strength, so some claycontent is desirable. Reducing ball size decreases the heat sensitivityof clay, perhaps allowing steam to escape more easily. The amount oflignin sulfonate also influences heat sensitivity, possibly because morelignin sulfonate produces a thicker coating that may limit escapingsteam, as with the clay. About 0.5 wt % lignin sulfonate was found toprovide the benefits described herein while further decreasing heatsensitivity observed at 1.0 wt %. Even so, 1.0 wt % produced a morerobust coating.

The drying occurs from the outside of the balls where heat is passed tothe water at the surface of the balls by convective heat transfer fromthe heated gases. As water evaporates as steam, it is replaced withwater from deeper within the ball. The new water passing to the surfaceof the ball is believed to bring with it additional lignin sulfonate.This process continues until all of the water has been dried from theballs.

The result is that on the order of half of all of the lignin sulfonateends up on the surface of the balls, forming a layer about 50 micronsthick. If less wicks to the surface, then the coating may be thinner,but more remains inside the ball. Apparently due to the wickingmechanism, the remainder of the lignin sulfonate may exhibit acontinuously decreasing concentration gradient of lignin sulfonate fromthe surface into the ball. As a result of the concentration gradient,observation of balls produced according to embodiments herein does notreveal a clear line of demarcation between the coating and the ballsurface. Instead, a smooth transition exists into the ball.

Another feature of the lignin sulfonate coating upon drying is that itis tough and adherent, undoubtedly responsible for the surprisingattractive process benefits described herein. It is further surprisingthat the surface coating of lignin sulfonate does not stick the ballstogether during the drying process, even though the balls are in contactand do not move relative to one another during the drying process. It ispossible that the evaporating water from the balls creates a vaporbarrier between balls that keeps them from sticking to one another asthe coating forms.

One beneficial result is that the strength of the green (as formed) and,especially, the dried ball, is surprisingly high. Their strengthsobtained usually exceed that of balls including instead an equal amountof phosphate clay. For example, the dry crush strength may be greaterthan 25 pounds (lbs). The drop strengths, which measure the number ofdrops from 18 inches a green or dried ball can sustain before breaking,is also enhanced. This means that fewer balls break at transfer pointsduring conveyor transfers, etc. in a commercial IHP. The strength mightbe attributed to having a coating covering the ball, a gradient oflignin sulfonate from the surface into the ball, or both.

The balls enter the kiln generally at about 120° C. in the IHP.Observation indicates that less carbon is burned out of the balls whenlignin sulfonate is used in the formulation. Apparently, as the ballsheat up in the back end of the kiln, the lignin sulfonate pyrolyzes,releasing hydrocarbon gases that pass to the kiln freeboard. These gasesform a gaseous layer that restricts the passage of P₄O₁₀ from the kilnfreeboard to the balls, where it may react with fluorapatite mineral inthe ball to form calcium pyrophosphate.

Without the lignin sulfonate coating, as the balls further heat up, theexcess phosphorus from the kiln freeboard may react with carbon in theball at a temperature lower than the main phosphate reduction reaction,removing carbon from the outer shell of the ball. When the carbon hasbeen removed in an area of the balls, then no reduction of phosphorusoccurs and the process yield goes down.

At about 650° C., the volatiles from the green petroleum coke componentof the ball begin to evolve and further blanket the balls as explainedin the Improved Hard Process section below. The carbon in the balls doesnot oxidize with water or CO₂ arriving from the freeboard at asignificant rate much below 650 to 700° C. Thus, until the pyrophosphatemechanism was understood during development of the IHP, it was notappreciated that lower temperature protection was of significance. Thelow temperature volatiles benefit from a larger kiln size as the bedsurface-to-volume ratio decreases. The flux of P₄O₁₀ coming from thefreeboard is related to bed surface area. The flux of low temperaturevolatiles from the lignin sulfonate increases as the volume of the bedincreases. Thus, as the kiln size is increased the blanketing of the bedis proportionately increased.

Note also that the hydrocarbon volatiles from the lignin sulfonate notonly blanket the bed, but also react with the P₄O₁₀, chemically reducingit as a first reaction step to phosphorus, carbon monoxide, andhydrogen, none of which can oxidize carbon. The amount of hydrocarbonvolatiles available for destroying incoming P₄O₁₀ is also directlyproportional to the size of the kiln. Thus, the lignin sulfonate can besignificant in protecting the balls entering the kiln from reaction withP₄O₁₀ in the back end of the kiln. The fixed carbon layer from thelignin sulfonate also further protects the fluorapatite from carbonburnout by forming a sacrificial layer that becomes proportionallygreater as the surface-to-volume ratio of the kiln decreases due to kilnsize scale up.

Another beneficial effect of the lignin sulfonate layer is due to itstough, adherent nature on the outside of the ball. It is during therapid heat-up region in the back of the kiln where much of the dustingoccurs. The dust passes to the freeboard of the kiln where it reactswith P₄O₁₀ to form a sticky calcium pyrophosphate material. It eitherpasses with the freeboard gasses into the hydrator, or passes to thewall of the kiln where it can adhere to the kiln wall, eventuallybuilding up enough to cause kiln ringing. The tough, adherent nature ofthe lignin sulfonate layer on the outside of the ball considerablyreduces the amount of dusting that occurs. Note that lower dusting alsoincreases the quality of acid that can be made in the IHP by reducingthe amount of dust components passing to the acid forming in thehydrator.

Another benefit of using lignin sulfonate is the observed small, butdistinct increased yield at low reaction temperature in the IHP in thelab pilot unit where carbon burnout is not operating. The observationmay partially result from having the lignin sulfonate carbon moreintimately in contact with the phosphate particles in the ball than arethe green petroleum coke particles.

Another benefit of using lignin sulfonate is its ease of use. If thesmall amount of lignin sulfonate (usually about 1.0% of the feed weight)is combined with enough water to form the moist mixture for balling(usually about 9% moisture) it forms a very liquid, low viscosity fluidthat can be added to the mixer with a pump and liquid sprays.

A further benefit is that it consumes waste from the sulfite paperproduction process that might not otherwise be profitably used.

Thus, in an embodiment, a phosphorous pentoxide producing methodincludes co-grinding a mixture containing phosphate ore, carbonaceousmaterial, and silica. A lignin sulfonate is homogeneously blended withthe co-ground mixture. Green feed agglomerates are formed from theblended lignin sulfonate and co-ground mixture, then dried, wicking someof the lignin sulfonate onto a surface of individual dryingagglomerates. The wicking forms a lignin sulfonate coating thereon. Thedried agglomerates contain phosphate ore particles, carbonaceousmaterial particles containing about 8 to 12 wt % volatiles, andsufficient silica particles for the dried agglomerates to exhibit acalcium plus magnesium-to-silica mole ratio of less than 0.5. At least80% of all particles exhibit a size less than 200 mesh.

The method includes forming a reducing kiln bed using the driedagglomerates below an oxidizing freeboard of a counter-current rotarykiln, the bed having a length. The bed is blanketed with volatilesevolving at least partly from the lignin sulfonate during bed heat upfrom about 120 to about 650° C. A sacrificial layer of fixed carbon isretained from the lignin sulfonate coating. The method includesvolatilizing at least a portion of the sacrificial layer during bed heatup from about 650 to about 1180° C., maintaining a bed temperature at orabove 1180° C. along a portion of the bed length, generating kiln offgas, and collecting phosphorous pentoxide from the kiln off gas. Thekiln discharges a residue containing processed agglomerates, less than10% of the agglomerates' phosphate input to the kiln remaining in theresidue.

By way of example, the homogeneous blend may further include water andthe drying may involve controlling a drying rate of the greenagglomerates. As the water evaporates from the agglomerates, the ligninsulfonate coating forms on individual dried agglomerates by wicking someof the lignin sulfonate onto the surface of individual dryingagglomerates without adhering the agglomerates together. The green feedagglomerates may be green feed balls exhibiting a greater than ¼ to lessthan ½ inch diameter prior to drying. The dried feed balls may exhibit adry crush strength greater than 25 pounds. The lignin sulfonate coatingmay have a thickness of about 10 to 40 micrometers and a remainder ofthe lignin sulfonate may exhibit a continuously decreasing concentrationgradient of lignin sulfonate from the surface into individualagglomerates. Individual agglomerates may contain a total of about 0.4to 2 wt % lignin sulfonate, including the coating.

In addition to the methods described herein, the embodiments may alsoinclude feed balls of previously unknown composition. Accordingly, in anembodiment, a phosphate ore feed ball includes a dry, homogeneousmixture of phosphate ore particles, carbonaceous material particlescontaining about 8 to 12 wt % volatiles, and sufficient silica particlesfor the ball to exhibit a calcium-to-silica mole ratio of less than 1.0,at least 80% of all particles exhibiting a size less than 200 mesh. Theball further includes a lignin sulfonate both inside the ball andcoating a surface of the ball. The coating has a thickness of about 1 to50 micrometers and a remainder of the lignin sulfonate exhibits acontinuously decreasing concentration gradient of lignin sulfonate fromthe surface into the ball. The ball contains a total of about 0.1 to 4wt % lignin sulfonate, including the coating. The ball exhibits a drycrush strength greater than 25 pounds.

By way of example, the carbonaceous material may include green petroleumcoke. The calcium plus magnesium-to-silica mole ratio may be less than0.5.

Given that some of the beneficial properties of the feed balls may bedifficult to quantify, the feed balls may be characterized by a processfor producing them, yielding the beneficial properties describedelsewhere herein. Accordingly, an embodiment provides phosphate ore feedballs produced by the process including co-grinding a mixture containingphosphate ore, carbonaceous material, and silica and homogeneouslyblending water and a liquid containing lignin sulfonate with theco-ground mixture. Green feed balls are formed from the blended water,lignin sulfonate, and co-ground mixture, the green balls containingabout 0.1 to 4 wt % lignin sulfonate.

The process includes drying the green balls by passing heated airthrough a multilayer bed of green balls and controlling a drying rate ofthe balls. As water evaporates from the balls, a lignin sulfonatecoating forms on individual dried balls by wicking some of the ligninsulfonate onto a surface of individual drying balls without adhering theballs together. A remainder of the lignin sulfonate in individual driedballs forms a continuously decreasing concentration gradient of ligninsulfonate from the surface into individual dried balls. The dried ballscontain phosphate ore particles, carbonaceous material particlescontaining about 8 to 12 wt % volatiles, and sufficient silica particlesfor the dried balls to exhibit a calcium-to-silica mole ratio of lessthan 1.0. At least 80% of all particles exhibit a size less than 200mesh.

By way of example, the green feed balls may exhibit a greater than ¼ toless than ½ inch diameter prior to drying. The dried feed balls mayexhibit a dry crush strength greater than 25 pounds. The carbonaceousmaterial may include green petroleum coke. The calcium plusmagnesium-to-silica mole ratio may be less than 0.5.

EXAMPLE

A dry mixture of 53.4 wt % phosphate pebble, 32.4 wt % silica tailingsfrom a Florida phosphate float plant, 5.2 wt % clay, and 7.2 wt % greenpetroleum coke was prepared and mixed. The mixture was dry ground to80%-200 mesh in a open circuit ball mill. The mixture was combined with9 wt % water in a mixer for 20 minutes to form blend A. A second mixtureinstead using 8 wt % water and 2 wt % calcium lignin sulfonate liquor(containing 50 wt % calcium lignin sulfonate) was prepared in the mixeras blend B. Both mixtures were formed into balls on a one meter ballingpan using a small admixture of water using water sprays and yieldingending moisture of about 11 wt %.

Both sets of green balls had a drop test value of >20, which means thatthey survived 20 drops from a height of 18 inches without breaking. Bothsets of balls were dried at a temperature of about 250° C. in a dryingoven overnight. The blend B balls appeared to have a black, leathery,tough coating that was difficult to scrape through with a thumbnailcompared to blend A balls, which lacked a coating and could be easilyscratched with a thumbnail. The dried blend A balls had a dry crushstrength of 18.6 lbs, while those with the lignin sulfonate added gave a42 lb dry crush strength. Levels over about 14 lbs indicate that theballs can be successfully processed in the kiln with acceptablebreakage. However, higher strengths indicate higher resistance todegradation in the kiln.

Due to the high turbulence in the batch kiln, tests were run with equalweights of bedding coke to reduce carbon burnout to a lower level. Forthese tests, 40 lbs of bedding coke (calcined petroleum coke) was addedto a batch kiln measuring 24 inches inside the refractory by 44 incheslong and preheated to 1350° C. A 40 lb charge of tests balls was thenadded and periodic grab samples allowed chemical analysis of the ballsto determine how much phosphorus had been released from the balls.

For the blend A balls the ultimate yield of phosphorus was 48% with anindicated carbon burnout rate of 2.8 pounds per hour per square foot.The amount of dusting from the balls was measured by determining howmuch calcium accumulated in a stainless steel spray tower with venturiscrubber train with water recycled from a small water tank. The calciummeasurement indicated a dust load of about 0.85% of the charge of ballsover a 95 minute test period. The high dust rate is believed due to thesmall size of the batch kiln and the high turbulence therein. The blendB balls processed under identical conditions gave a phosphorus yield of72%, a carbon burnout rate of 1.65 pounds per hour per square foot, anda dusting rate of 0.53%. All of these values are higher than wouldusually occur in a continuous countercurrent pilot plant test of theHard process carried out in a 2 ft 9 in ID by 30 ft long kiln. In thecontinuous kiln the turbulence is much less and the carbon burnoutreduced by an order of magnitude.

Improved Hard Process

Through research and development of the KPA process, observation, study,and insight resulted in several process improvements that overcomedeficiencies of the processes described in Hard and other referencesmentioned in the Background section above. Various embodiments of theIHP provide for high yields of phosphoric acid, efficient use ofcarbonaceous raw material requirements, high kiln throughputs, longrefractory life, limited sinoite (Si₂N₂O) formation, oxidizing off gas,and the capability of limiting sodium, potassium, fluorine, and sulfurloss from the kiln solids.

Efforts have confirmed that the yield limitations present in the Hardcontinuous pilot plant resulted primarily from burn out of carbon in ashell that formed on the outside portion of the agglomerate as it heatedup to greater than 1180° C., at which temperature the phosphate orereduction began. The carbon burned out (oxidized) from the outsidesurface of the ball by reaction with oxidizing gases from the kilnfreeboard (e.g., CO₂, P₄O₁₀, and H₂O). Assuming an agglomerate was inthe form of a 1.0 centimeter ball and 30 weight % (wt %) of the carbonwas burned out, then a white shell about 800 micrometers thick formed onthe outside of the ball. The phosphate ore in the shell was notchemically reduced, even after reaching reaction temperature as therewas no carbon therein to chemically reduce it, and yield was limited.

Also, sodium, potassium, sulfur and fluorine were largely volatilizedfrom the shell, but were retained in the core containing residualcarbon. Thus, limiting carbon burnout may reduce loss of sodium,potassium, sulfur and fluorine from the bed.

The Megy IHP patents teach that after phosphate reduction begins and thephosphorus metal vapor begins to form in the core of an agglomerate,phosphorous metal may encounter oxidizing gases either in the oxidizedshell of the ball or in the bed between the balls. The P₄ emerging fromthe core may oxidize into P₄O₁₀, which may back react with the unreactedphosphate ore in the shell to form calcium pyrophosphate. Calciumpyrophosphate has a melting point of 973° C. (which is below thereduction temperature) above which it forms calcium metaphosphate.Calcium metaphosphate is in equilibrium with the vapor pressure of P₄O₁₀and maintains a composition near to that of calcium pyrophosphateprovided the P₄O₁₀ vapor pressure is greater than 1 millimeter ofmercury. Such formation of calcium metaphosphate by back reaction ofP₄O₁₀ with fluorapatite likely further limited yield in the Hard processbeyond yield loss directly associated with carbon burnout.

However, efforts taken to limit carbon burnout may also decrease thetransfer of P₄O₁₀ to the bed and, therefore, decrease P₄O₁₀ reactionwith ore and formation of calcium metaphosphate. Thus, it becomes doublysignificant to reduce carbon burnout from the bed in order to limit theindicated back reaction, further increasing yield and also retainingmore sodium, potassium, sulfur and fluorine otherwise lost from the bed.

If the phosphorus metal vapor and associated carbon monoxide from thereduction reaction evolve fast enough, then they sweep the oxidizinggases away from the bed and limit further oxidation of the carbon fromthe bed. The vigor of the evolution is directly related to the bedsurface-to-volume ratio. In sufficiently large kilns, the carbon burnoutis effectively stopped during the period when the phosphorus reductionreaction is occurring. Once the reduction reaction occurs, the solidreaction product includes an amorphous (glassy) calcium metasilicate(CaSiO₃), which encapsulates any remaining carbon, protecting it fromfurther oxidation.

Since loss of carbon from the ball by carbon burnout may lead to areduction in yield, limiting the carbon burnout to low levels mayprovide high yields. Control of carbon burnout may be achieved primarilyby implementing one or more of the following measures, which areadditive in their effect in reducing carbon burnout:

1) Reducing surface-to-volume ratio of the kiln bed.

2) Using a control system which includes air ports down the length ofthe kiln to introduce over bed air at a controlled rate to provide for arapid heating of the kiln feed to the temperature where the phosphatereduction reaction proceeds (i.e. greater than 1180° C.) and maintainingthe reduction temperature over most of the length of the kiln.

3) Using appropriate kiln design to reduce turbulence in the region ofthe kiln where the kiln solids heat up from 700 to 1180° C., where thephosphate reduction reaction begins.

4) Using uncalcined carbonaceous raw material to provide an off gas ofhydrocarbons during the heating of the kiln feed to reaction temperatureto blanket the bed in reducing gasses and to react with oxidizinggasses.

5) Using finely ground and intimately mixed carbonaceous material in thefeed agglomerates to lower the temperature of full phosphate reductionat commercial feed rates to about 1180° C.

6) Adding a small, but sufficient, amount of uncalcined carbonaceous rawmaterial together with the kiln feed agglomerates as bedding topreferentially react with incoming oxidizing gases, particularlyphosphorus oxide, from the freeboard.

According to the IHP, a phosphorous pentoxide producing method includesforming a kiln bed using feed agglomerates in a counter-current rotarykiln, the bed having a length and the agglomerates containing phosphateore particles, carbonaceous material particles, and sufficient silicaparticles for the agglomerates to exhibit a calcium-to-silica mole ratioof less than 1.0. The method includes maintaining a bed temperature ator above 1180° C. along at least 50% of the bed length without exceeding1380° C. along the entire bed length. Kiln off gas is generated andphosphorous pentoxide is collected from the kiln off gas, the kilndischarging a residue containing processed agglomerates and less than10% of the agglomerates' phosphate input to the kiln remaining in theresidue as phosphate.

The Megy IHP patents teach that reducing bed surface-to-volume ratioperhaps provides the most significant benefit, particularly incombination with the use of uncalcined petroleum coke or coal as thecarbonaceous raw material source in the agglomerates. To process a givenbed volume, a stoichiometric amount of oxygen is provided to generatesufficient heat in the off gas to run the process. Supplying the oxygenfrom air moves a specific volume of kiln freeboard gases countercurrentto the kiln solids.

In scaling up, commercially-sized kilns normally exceed a minimumdiameter above which kiln shell heat losses become minor. Thecross-sectional area of the bed and freeboard both scale up with thesquare of the diameter of the kiln. Since the length to diameter ratioof the kiln may be similar regardless of the size of the kiln, the bedvolume scales up as the cube of the kiln diameter. The stoichiometricamount of oxygen and the corresponding specific volume of freeboardgases from the air that provides the oxygen are roughly proportional tothe bed volume. The cross-sectional area of the freeboard is directlyproportional to kiln diameter squared, so the velocity of the gasflowing countercurrent to the kiln solids increases with increasing kilndiameter at a given temperature of operation.

The gas velocity relative to the kiln solids increases with the kilndiameter and the turbulence increases somewhat, but more slowly than indirect proportion to kiln diameter. However, the transfer of oxidizinggases from the freeboard into the bed is reduced, both due to the slowerincrease in mass transfer associated with the increase in kiln gasvelocity and due to the multiple and independent protections of theblanketing and oxidizing gas interceptions taught herein. Thesemechanisms multiply the effect of kiln scale up and rapidly diminish thecarbon burnout rate. The kiln may include dams to better control thedepth of the solids over the kiln length. The dams may have a long,tapered approach to limit the turbulence that may otherwise result fromtheir use.

The surface area available for heat transfer from the freeboard gassesto the kiln solids in the bed is directly proportional to the length ofthe kiln. If stoichiometric air is used to oxidize the phosphorus,carbon monoxide and hydrocarbons from the bed, then the total flow ofgas in the kiln increases for an increased bed volume. Thus, the gasvelocity in the kiln is directly proportional to the kiln length exceptthat the kiln shell losses decrease with kiln length.

Long kilns are used in heat transfer processes to maximize heatutilization in processes like wet process cement production etc. In aKPA process, a stoichiometric amount of carbon is used to chemicallyreduce the phosphorus from the bed. The reduction produces a specificamount of combustibles from the bed which, when oxidized with air, givea modest excess of heat in the kiln. The excess heat gives a kiln offgas temperature higher than would be required. A direct result is that akiln suitable for the KPA process may be shorter rather than longer. Theshorter kiln presents a condition that also benefits carbon burnout asthe freeboard gas velocity is reduced. Thus, in the KPA process, thecarbon burnout rate increases in longer kilns as the freeboard gasvelocity increases.

Evaluation of three data sources provided the carbon burnout in batchand commercial kilns. As the first source, sufficient data is availablein Folmo et al., Ilmenite Direct Reduction Project in Norway using theGRATE-CAR™ Process, AIME Conference, 1992, supplemented by informationfrom Allis Chalmers engineers that started up the facility. The datapertains to carbon burnout rate in a 19 feet diameter by 232 feet longchemically reducing kiln processing ilmenite ore in Tyssendal, Norwaysince 1986 to produce metallic iron and a higher value TiO₂ residue. Thekiln solids were held at about 1170° C. and the carbon burnout was about1.1 pound per hour per square foot. Although not enough data is given inthe Folmo article to estimate the freeboard velocity, the kiln was a bitelongated, which probably resulted in a freeboard gas flow rate ofperhaps 25 feet per second, as is common in kilns of its shape.

As the second source, in the Leder article discussed in the Backgroundsection above, the temperature of the kiln bed averaged about 1200° C.over the length where the carbon burnout was determined to average inthe range of 0.2 to 0.3 pounds per hour per square foot. The gasvelocity in this much shorter kiln averaged about 3.3 feet per second.

As the third source, a 13.5 inch batch kiln used in IHP Example 3described below was determined carefully to produce carbon burnout ofabout 2 pounds per hour per square foot during the time the solids wereheating to reaction temperature.

The first observation that can be made from the above data involves theburn out of carbon as a first approximation being directly related tothe surface area of the kiln bed, which in turn is directly related tothe diameter of the kiln and the length of the kiln. In the twocontinuous kiln systems above, the amount of surface in the kiln systemschanges by a factor of 55, but the carbon burnout only by a factor of3-4. Most of the higher burnout rate in the larger kiln is due to theincreased flow rate of the freeboard gasses. Also, in the Tyssendal kilnwith bedding coke, the burnout tends to occur over the entire length ofthe kiln whereas, in the IHP Example 3 batch kiln, the carbon burns outmostly during the charge heat up period.

The carbon burnout rate in the batch kiln is a factor of 7-10 higherthan the continuous kiln even though the “apparent” freeboard gasvelocity is only 4.4 feet per second. The batch kiln is closed with aperpendicular lid with a 9 inch hole in it. In fact all of the kilngases exit through the relatively small hole at an exit velocity ofabout 30 feet per second. The gas flow through the relatively small holeentrains other gas in the kiln that diverges toward the bed as it hitsthe lid, setting up a strong recycle gas flow. When this gas flow hitsthe bed at the lower portion of the lid, it impinges directly on thesurface of the bed and the momentum drives the oxidizing gases directlyinto the bed, resulting in enhanced carbon burnout. The high turbulencein the batch kiln causes high carbon burnout and limits high yield inreduction processes in the batch kiln. The batch kiln represents thestandard tool commonly used in the development of kiln processes. Theenhanced burnout associated with the turbulence in the batch kilnapparently led researchers to underestimate the prospects for chemicallyreducing bed processes.

In a continuous kiln operated as previously known, the feed agglomeratesenter at about 125° C., and are well away from the inlet end of the kilnby the time they reach the temperature of 700° C. where the carbon burnsout of the bed fast enough to be of concern to the process. At thatpoint, the gases in the kiln flow smoothly over the kiln solids withminimum turbulence. In embodiments of the IHP, increased turbulencemight be reduced by tapering a dam at the ends of the kiln rather thanusing an abrupt dam to retain solids in the bed.

The amount of carbon present in the kiln bed relates to the volume ofthe bed, which relates to the square of the diameter of the kiln perunit length. The carbon burnout is proportional to the surface area,which relates to the first power of the diameter of the kiln per unitlength. Therefore, the fraction of carbon that burns out of the kiln beddirectly relates to the inverse of the diameter of the kiln. As shown inIHP Example 3 below, the magnitude of the rate of carbon burnout is suchthat in a small batch kiln (e.g., 13.5 inch diameter batch kiln) it isdifficult to have any carbon left in the kiln solids after heating themto reaction temperature. In large, continuous kilns a much greaterfraction of the carbon remains in the kiln solids.

Another factor that directly affects the amount of carbon burnout is howlong the kiln solids are exposed to oxidizing freeboard gases as theyheat up to reaction temperature. In the KPA process, rapidly heating thefeed agglomerates to reaction temperature involves a means of oxidizingpart of the phosphorus and carbon monoxide off gases from the phosphatereduction reaction down the length of the kiln using the kiln controlsystem described below for FIG. 1 and/or IHP Example 5.

In a 6 foot diameter continuous pilot plant kiln (72 feet long) designedfor low turbulence, a kiln system with the ratio of bedsurface-to-volume multiplied by the bed heat up time of less than 50minutes-ft²/ft³ may obtain yields greater than 92% with high throughput,low impurity volatilization, extended refractory life, and eliminationof the competing sinoite reaction described below. As explained below infurther detail, larger diameter kilns of greater than 12 feet operatingunder otherwise identical conditions may provide greater yields, forexample, greater than 95%.

In combination with the use of uncalcined carbonaceous material in thefeed agglomerates, surface/volume ratio in the kiln bed provides evenmore benefit. Uncalcined carbonaceous material decomposes when heatingto form gaseous hydrocarbons (mostly methane) and a stable carbonresidue known in the industry as “fixed carbon,” which does notvolatilize at process temperatures and is the reactant used tochemically reduce the phosphate ore. The volatiles in carbonaceouscomponents of phosphate ore and coal begin to evolve as soon as the feedagglomerates are heated above room temperature. Green petroleum coke,having already been subjected to coking temperatures, may begin evolvinghydrocarbons at about 600° C. Most of the carbonaceous materials finishevolving their hydrocarbons by about 950° C.

During bed heat up, oxidation of fixed carbon by air at a significantrate often does not occur until reaching temperatures above about 700°C., well after the volatiles from green petroleum coke in kiln solidsbegin to exert their protective effect at about 600° C. Even so,transfer of P₄O₁₀ to the bed may occur from the time the solids enterthe kiln until they reach reaction temperature. Thus, the time it takesto bring the kiln solids from kiln inlet to 1180° C. defines “heat uptime” in the context of the present document, unless otherwiseindicated. However, it is clear that the availability of uncalcinedcarbonaceous material in the kiln solids makes the process somewhat moretolerant to a slower kiln solids heat up rate.

During the vulnerable period when feed agglomerates are being heated toreaction temperature, observation indicates that volatilization ofgaseous hydrocarbons provides a gas sweep, significantly limiting thetransport of oxidizing species from the kiln freeboard to the kiln bed.The extent of the hydrocarbon gas sweep is directly proportional to thediameter of the kiln. That is, the amount of volatiles present in thebed is directly proportional to the volume of the bed (i.e. the squareof the diameter of the kiln per unit of length) while the area that thevolatiles come out of the bed is proportional to the area of the bed(i.e. the diameter of the kiln per unit of length). Thus, the flux ofhydrocarbons from the surface of the bed is directly proportional to thediameter of the kiln.

However, the amount of volatiles in green petroleum coke is on the orderof ten times less than the amount of phosphorus metal and carbonmonoxide formed after a given volume of agglomerates reaches 1180° C.Therefore, the effectiveness in sweeping away oxidizing gasses from thefreeboard increases with the scale up of the kiln to larger diameters,making commercial-scale operations more important to realize thisprotection.

It is also important not to allow the volatiles in the petroleum coke toget too high since they provide heat to a kiln system with an oxidizingoff gas and, if too high, then they make operating an oxidizing kilnmore difficult. In addition, the volatiles add hydrogen to the kiln,ultimately producing water that limits the strength of phosphoric acidthat can be made from the process. However, green petroleum coke withless than 12 wt % volatiles may be suitable for use in the KPA process.Reduction in carbon burnout with increasing kiln diameter due to thesweep effect of the volatile gases multiplies the reduction in carbonburnout with increasing kiln diameter due to the surface area limitationof oxidizing gas transport to the bed, thus giving compounded loweringof carbon burnout with kiln scale up.

The evolving hydrocarbons from the kiln bed may have another directeffect on the reduction of carbon burnout from the kiln bed. Thevolatile hydrocarbons may react with the oxidizing gases from thefreeboard in a way that creates a product which is not reducible by thecarbon in the kiln bed. Consider the following reaction:

CH₄(from bed)+CO₂(from freeboard)>CO+2H₂

Neither of the reaction products is reducible by the carbon in the bed.As a result, the hydrocarbon not only sweeps the oxidizing gasses awayfrom the bed, but also intercepts the remaining oxidizing gases andtransforms them into nonoxidizing gases. Note that this effect is alsodirectly related to the surface/volume ratio in the kiln bed andmultiplies the reduction in carbon burnout fraction by the twomechanisms discussed above.

Even though the use of uncalcined carbonaceous material evolvesvolatiles effective in substantially reducing carbon burnout in largeenough kilns, its evolution from the bed is largely complete at 950° C.,leaving the bed largely unprotected between 950 and 1180° C., where thephosphate ore reduction reaction proceeds at a reasonable rate. In orderto minimize the carbon burnout during that interval, incoming kilnsolids may be heated to reaction temperature as rapidly as possible.

Special provisions may be made to rapidly heat up the kiln solids. TheKPA process is unusual in that essentially all of the fuel required toheat the kiln and feed agglomerates and to provide the heat for theendothermic phosphate reduction reaction is generated as a result of thereduction reaction itself. The reaction forms phosphorus metal andcarbon monoxide off gases according to the reaction:

Ca₁₀(PO₄)₆F₂+9SiO₂+15C> 3/2P₄(gas)+15CO(gas)+9CaSiO₃+CaF₂

The combustion of the carbon monoxide and phosphorus metal from thereduction reaction, the burning of coke volatiles, and the oxidation ofsome of the carbon from the bed generates all the heat necessary for theKPA process. Nevertheless, the heat must be distributed down the kiln tobe effectively used. The hot combustion gases that result from theburning of phosphorus and carbon monoxide do not have sufficient latentheat to supply the endothermic needs of the phosphate reductionreaction. Therefore, previous attempts by Hard, Megy, etc. to run theKPA process by passing air entirely into the burner end of the kilnresulted in runaway temperatures over a small length of the kiln nearthe burner.

Such a runaway reaction occurred in the continuous kiln pilot plant ofthe Leder article discussed in the Background section above and not onlylimited its performance, but forced it to be operated at high silicafeed formulations, chemically reducing off gas, and a high temperaturepeak near the burner end. Together with modest yields, the processingrestrictions resulted in lowered commercial prospects for the process.One means to spread the heat generated by the burning of the phosphorusmetal gas and carbon monoxide down the kiln is to limit the amount ofoxygen present at the site of the reaction near the burner. Limitingoxygen leaves phosphorus and carbon monoxide in the freeboard which isthen burned with cold air from over bed air ports as needed to supply aneven temperature down the kiln.

Ported kilns are not common, but have been used previously in severalcommercial kiln processes to supply over bed air for down kilncombustion or place air or other gases under the bed as a reactant.Using a ported kiln to supply controlled over bed air to the KPA processallows the temperature to be held at an optimum temperature profile forforwarding the phosphate reduction reaction over most of the kiln lengthand to bring the incoming kiln solids rapidly up to reactiontemperature. By reducing the time the kiln feed agglomerates spend inreaching reaction temperature the carbon burnout may be reduced. Asdiscussed below in relation to FIG. 1, a control system may beimplemented to achieve the desired effects of over bed air addition.

Several other benefits accrue to the establishment of a long reactionzone at a just sufficient minimum kiln bed temperature, including: muchincreased processing rates due to the longer reaction times for heattransfer, longer refractory life, reducing clinker formation andrefractory-kiln solids interactions, eliminating sinoite formation, andminimizing the loss of sulfur, sodium, potassium, and fluorine from thebed.

The Megy IHP patents teach that the sinoite reaction occurs during theoperation of the KPA process if run at higher temperatures. The reactionof nitrogen from the freeboard and silica with fixed carbon in thereducing kiln bed competes with the phosphate reduction reaction above1310° C. Between 1180° C. and 1310° C. the phosphorus reaction occurs,but not the sinoite reaction. If the KPA kiln residue is heated to over1310° C. after the phosphate ore has been chemically reduced, then thesinoite reaction proceeds, using up any excess carbon. When the carbonis exhausted from the kiln solids residual, it becomes stickier and moredifficult to handle. The sinoite does not otherwise help the KPAprocess, so it is useful to get maximum throughput and discharge thekiln solids from the kiln as soon as the phosphate ore has beenchemically reduced.

One additional method of limiting carbon burnout in the KPA kilnincludes adding a small amount of loose, small sized pieces of acarbonaceous material together with the raw material feed agglomeratesto the KPA kiln. If the kiln has a small enough bed surface/volume ratioand a fast enough feed heat up rate, then small amounts of carbonaceousmaterial at about 1-2% of the weight of the feed agglomerates aresufficient to provide sacrificial carbon burnout throughout the feedagglomerates while heating them to reaction temperature. Such additionsof sacrificial carbon outside the feed agglomerates are more effectivethan adding more carbon to the raw material mixture that make up theagglomerates.

An advantage exists in using uncalcined carbonaceous material since thevolatiles therefrom contribute to the volatiles sweep effect and to thereaction with the incoming oxygen flux from the freeboard, decreasingP₄O₁₀ condensation and calcium metaphosphate formation. It isparticularly advantageous to use carbonaceous material that evolves somevolatiles between 100° C. and 650° C., which limits P₄O₁₀ transfer fromthe kiln freeboard to the cold raw material agglomerates entering thekiln. The P₄O₁₀ otherwise condenses on the cold feed balls at less than485° C. and can react with fluorapatite (Ca₁₀(PO₄)₆F₂) to form calciumpyrophosphate. Additionally, P₄O₁₀ may react with carbon at about 800°C. Enrichment of feed balls with phosphorous in such manner essentiallyrecycles phosphorous from the back end of the kiln to the front end.Such enrichment accentuates carbon burnout by placing an increaseddemand on carbon to re-reduce phosphorous in the condensed P₄O₁₀ and/orcalcium pyrophosphate. Although the slow mass transfer from the kilnfreeboard to the bed limits the extent of this occurrence, particularlyin larger kilns, it can be further mitigated by the low temperaturevolatilization of organics from the carbonaceous material.

If the phosphate ore has a suitable organic level it is useful for thispurpose. Coal also has low temperature volatile organics that is usefulfor this purpose. The carbonaceous material may be more effective ifsized to small pieces that present maximum surface area to the incomingoxygen flux. The particle size should not be so small that it entrainsas dust into the freeboard gases of the kiln. Carbonaceous material inthe ¼ inch to ½ inch size range has been found to be effective.

Fine grinding of the carbonaceous material and its effectivedistribution in the raw material agglomerates may be important to theprocess. Jacob et al., Reduction of Tricalcium Phosphate by Carbon,I&EC, Vol. 21, No. 11, 1929, 1126-1132 and Mu et al., Thermodynamics andKinetics of the Mechanism of Reduction of Phosphate Ores by Carbon, Met.Trans. B, 17B, 1986, 861-868 discuss that phosphate reduction withcarbon is a first order reaction, but neither describe the effect ofcarbon surface area on the rate of reaction or the effect of intimatemixing. Observation indicates the rate controlling step in the KPAprocess is the diffusion of phosphorus and silica molecules across agrowing calcium monosilicate layer around each carbon particle. As thelayer thickens, the reaction slows down, giving the known first orderreaction rate behavior.

Grinding the carbon finer not only gives a higher surface area for agiven weight of carbon, but also decreases the distance between thecarbon particles. Thus, the diffusion path to the carbon particles isshortened for the phosphate and silica. The Megy IHP patents teach thatthe carbon particle diameter shrinks as it reacts, resulting in a voidbetween the carbon and the phosphate ore/silica that may add extraresistance to phosphate reduction when the carbon particle size isgreater than 200 mesh. With such larger sized carbon particles, a slowerreaction rate exists due to a lower specific surface area. Also, theadded distance between particles results in a logarithmic increase inthe time it takes at a given temperature to consume all the phosphatebetween the particles.

Intimate mixing of the carbon ensures realization of the advantageousdiffusion characteristics and may be efficiently accomplished byco-grinding the raw materials. Otherwise, carbon particles, even thoughsmaller, may clump together and provide less benefit since a clump ofsmall particles functions like a single large particle. Maintaining ahigher reaction temperature can increase the phosphate/silica diffusionrate to overcome a lack of mixing or a larger carbon particle size.However, the higher reaction temperature may reduce kiln throughput,increase volatilization of bed impurities, increase sinoite productionand loss of phosphate yield, increase carbon burnout, lead to reducedrefractory life, and create other problems. Observation indicates thatopen-circuit ground green petroleum coke, 80% of which exhibits a sizeless than 200 mesh, providing an amount of fixed carbon 1.3 times atheoretical carbon requirement for reduction of all phosphate in the orecan carry out over 95% phosphate reduction in less than two hours whenintimately mixed in the agglomerates. However, over 95% phosphatereduction takes over six hours if −140/+200 mesh coke is used instead.“Open circuit” grinding refers to batch grinding until a charge ofmaterial meets specification as opposed to “closed circuit” wherematerial meeting specification is continuously removed and othermaterial remains for further grinding. Open circuit grinding thusgenerates more fines.

According to another embodiment of the IHP, a phosphorous pentoxideproducing method includes forming a kiln bed using feed agglomerates ina counter-current rotary kiln, the bed having a length and theagglomerates containing phosphate ore particles, carbonaceous materialparticles, and sufficient silica particles for the agglomerates toexhibit a calcium-to-silica mole ratio of less than 1.0. Individualagglomerates substantially exhibit a same elemental composition, a samecalcium-to-silica mole ratio, and a same proportion of excess fixedcarbon compared to a theoretical carbon requirement for reduction of allphosphate in the ore. Co-grinding and other concepts described hereinmay provide such agglomerates. The method includes maintaining a bedtemperature at or above 1180° C. along a portion of the bed length andestablishing a bed surface-to-volume ratio multiplied by a time for bedheat up to 1180° C. of less than 50 minutes-ft²/ft³ to obtain a yield ofgreater than 90%.

By way of example, the bed temperature may be maintained at or above1260° C. along at least 50% of the bed length. Maintaining the bedtemperature may include not exceeding 1380° C. along the entire bedlength. The method may further include blanketing the bed with volatilesevolving from the carbonaceous material during bed heat up from about600 to about 950° C. Also, the kiln bed may include, initially, fromabout 1 to about 2 wt % bedding coke or coal. Such bedding may blanketthe bed with volatiles evolving from the bedding coke or coal during bedheat up from about 100 to about 650° C.

Over bed air and/or oxygen may be added through a plurality of portsalong the bed length, the ports being of sufficient location, number,and throughput to decrease a time for bed heat up to 1180° C. comparedto otherwise identical processing without the over bed air and/oroxygen. Control systems may provide for monitoring of temperature at aninner surface of the kiln at a plurality of points along the bed length,the inner surface contacting the bed, and also monitoring oxygen andcarbon monoxide content of the kiln off gas. Further monitoring ofcarbon dioxide content may provide a check against the values obtainedfor oxygen and carbon monoxide content.

The Lapple and Megy patents introduced in the Background section specifya calcium-to-silica mole ratio greater than 1.0. However, observationindicates that a eutectic composition between the calcium metasilicateproduct from the reduction reaction and unreacted fluorapatite arisespart way through the reduction step in such processes and results inmelting problems in attempts to carry out such processes in rotarykilns. It is believed that all attempts to carry out the KPA processwith calcium-to-silica mole ratios greater than 1.0 cannot be successfulunless implementing very high calcium-to-silica ratios, as taught in thePark patent. Consequently, the embodiments of the IHP herein areadvantageously limited to a ratio less than 1.0.

FIG. 1 shows a block diagram demonstrating process flow for the KPAprocess. A range of phosphate ores may be processed by the KPA processby adjusting the raw material composition to accommodate the impuritylevel therein. For example, if the raw material mixture contains lessthan about 1.5 wt % Al₂O₃, then silica admixtures may be processed withcalcium-to-silica mole ratios approaching 1.0. On the other hand, ifores up to about 2 wt % Al₂O₃ are processed, then the calcium-to-silicaratio may be less than 0.5 to reduce melting problems in the kiln andprovide enough temperature window for commercial operation. No needexists to remove fines from the ore. Some fines are beneficial informing strong agglomerates for the KPA process. If washed ore is usedfor impurity control or to reduce moisture in the feed ore, thenbentonite or other clay additions to the raw material mixture help tomake sufficiently strong balls.

The amount of silica added also affects the overall heat balance of thekiln. Greater than stoichiometric amounts of fixed carbon are used tofully chemically reduce the phosphate ore in the KPA process. Thephosphate ore reduction produces a stoichiometric amount of phosphorusmetal vapor and carbon monoxide fuel. Thus, the KPA process tends todevelop more than enough heat to be utilized in the kiln with anoxidizing off gas. Additional silica in the mixture uses up some of theheat albeit at the expense of additional grinding and handling costs.

Phosphate ore deposits, as mined, frequently have sufficient silicacontent to meet the requirements of the KPA process. Sometimes a mixtureof the washed and unwashed ore is used to provide sufficient clays forbinder, but some clay removal may occur to meet the desired aluminumlevel for a given silica addition level. Potassium and sodium level inthe ore can be a problem if at unusually high levels. Iron impuritiesmay reduce yield by combining with phosphorus to form a ferrophosphorusalloy, which stays in the kiln residue. At levels normally found in theUS phosphate ore deposits, iron reduces yield to phosphorus oxide in theoff gas by 1.5-3%.

In some cases it may be economically advantageous to use phosphate orewith pebble phosphate rock removed for other uses and the remainingsilica, phosphate rock mixture beneficiated by fatty acid floatation toadjust silica level. High moisture levels in the as mined phosphate oremixture that are expensive to dry may affect raw material selection forthe process. Clay added to the process usually contains high associatedmoisture levels, but usually has a particle size distribution largelyunder 200 mesh. Thus, the clay need not be ground and wet clay can beadded to the other dry ground raw materials, also supplying the waterrequired in the agglomeration or balling operation. Such measure reducesthe overall drying requirements when wet clay is the binder.

The selection of the phosphate ore to use in the KPA process may be madeto minimize its cost. If one mine serves both KPA and dihydratephosphoric acid production plants, then additional options are availablein selecting phosphate ore from one of the partially beneficiated steamsas a component of the KPA raw material mixture.

In dry ore deposits, the phosphate ore feed to the KPA process may beas-mined phosphate ore matrix. Part of the matrix may warrant washing ifthe Al₂O₃ concentration is too high and silica additions may be desiredif the ore does not contain enough silica.

Sufficient carbonaceous material may be added to the phosphate ore toresult in fixed carbon content in the raw material mixture of about 1.3times the theoretical carbon required to chemically reduce the phosphatecontained in the raw material mixture. The volatile content of thecarbonaceous material may be limited to less than 12 wt % to achieve afavorable heat balance in the KPA process kiln (maintaining an oxidizingoff gas) and to limit hydrogen addition to the raw materials. The levelof hydrogen input to the kiln limits the maximum concentration ofphosphoric acid that can be made in the KPA process. Green petroleumcoke may be used as the carbonaceous material with volatiles in the 8-10wt % range, where they are high enough to offer protection againstcarbon burnout in larger kilns, but not so high that hydrogen additionsinterfere with producing 72-76% phosphoric acid from the KPA process.

High sulfur, green petroleum coke may be used in the KPA process sincemost of the sulfur stays with the kiln residue in an optimally runprocess. With power production as the main alternative use for petroleumcoke, high sulfur petroleum coke is generally cheaper than low sulfurpetroleum coke due to the expense of removing sulfur from the stack gasof power plants. While coal may be equally effective for such purposesin the KPA process, only coal with unusually low volatiles has lowenough hydrogen levels to curtail heat balance problems and limitationof the concentration of the phosphoric acid product. Alternatively, thecoal can be calcined, but then the cost of additional processing becomesa factor.

An example of a suitable raw material feed mixture for a favorabledeposit in the US state of Idaho is one part as-mined raw ore matrixwhich has been processed to remove material larger than 3 inches and0.09 parts green petroleum coke sized 3 inches and down. To supply a19-foot diameter (inside the bricks) rotary kiln to make 200,000 shorttons per year (P₂O₅ basis) in the form of super phosphoric acid (SPA)requires about 230,000 pounds/hour (lbs/hr) of raw matrix ore and 23,000lbs/hr green petroleum coke.

In process flow of FIG. 1, the raw materials are analyzed for moistureand metered together to form a raw material mixture then dried to 1.5 wt% moisture in a rotary drier using waste heat from the rotary coolerused later in the process. The dried ore mixture is sized to about ⅜inch in a crushing/screening circuit and passed into an open circuit dryball mill for grinding to 80%-200 mesh. Next, 9 wt % moisture is addedto the co-ground raw material mixture and the result formed into ballsin a balling drum/roller screen circuit with the addition of 3 wt %additional water in the balling drum sprays to control the ball growth.The balls exhibit a greater than ¼ to less than ½ inch (+¼/−½) diameter.

Thereafter, the wet green balls are layered on a perforated stainlesssteel belt and dried in a low temperature belt drier with waste heatfrom the kiln spent solids cooler later in the process. Further, 1 wt %(i.e., 2,500 lbs/hr) bituminous coal sized to +¼/−½ inch is added to thewet feed balls as they pass into the continuous belt dryer. Theagglomerates/bituminous coal feed leaves the drier at 125° C. and passesinto a 19 foot diameter (inside the bricks) rotary kiln.

The kiln of FIG. 1 is a ported kiln equipped with a circumferentialchannel with a moving seal that passes air from blowers outside the kilnthrough headers with volume control valves and ports to the inside ofthe kiln wall where the air is distributed along a channel in the brickand then to the freeboard of the kiln through slits in the brick. Shutoff valves on each of the headers are turned on and off during eachrotation to pass air into the freeboard only when the ports are abovethe kiln bed. The Folmo article referenced above describes such a kiln.

However, the kiln of FIG. 1 additionally includes a temperature and offgas control system. The kiln is fitted with on board thermocouples every6 feet which measure the temperature of the brick about ⅜ inch below thehot surface of the brick. The temperature seen by these thermocouplescan be related to the brick surface temperature as measured by anoptical pyrometer when the kiln is empty. When the KPA process isoperating, the white, opaque phosphorus burning flame obscures the viewof an optical pyrometer, which is the common way to determine internalkiln temperatures in other processes. Even so, the temperature dataobtained from the thermocouples may be relied upon to control airaddition through the ports and, thus, temperature profile along the kilnlength.

The kiln feed agglomerates are rapidly heated to 1180° C. by the hotcounter currently flowing freeboard gasses and maintained above thattemperature over most of the kiln length by adjusting the addition ofair along its length to achieve the corresponding onboard brickthermocouple readings. The feed end of the kiln includes a tapered damto reduce the turbulence associated with a dam commonly used in thesekilns to retain solids. Reduction in turbulence at the feed end achievesgreater benefit than at the burner end since the reaction is completedat the burner end. A small flame is maintained at the burner tostabilize the front of the kiln and achieve overall heat balance.

An off gas analyzer capable of analyzing CO, CO₂, and O₂ is included andcontinuously monitored to ensure the kiln maintains a slightly oxidizingoff gas. The above air addition control system may maintain the desiredkiln temperature profile. However, the off gas analyzer may be used tomonitor whether other process changes might be warranted, in keepingwith the discussion herein, so that the air added is also sufficient tooxidize all of the P₄ and CO, with O₂ in slight excess.

The hot kiln solids in FIG. 1 pass out of the kiln past an isolationcurtain into a rotary cooler. Heat is recovered from the hot balls intoan air stream, which is used to dry the combined raw materials and thegreen balls from the balling drum. The cooled balls pass out of thecooler, through an isolation valve to a conveyor belt that removes thespent balls to a disposal site.

The hot gases containing the P₄O₁₀ product exiting the FIG. 1 kiln atabout 1100° C. pass through a dust cyclone and then go to a spray towerthat contacts the off gas with recirculated, cooled strong phosphoricacid that absorbs about one half of the P₄O₁₀ therein and cools the offgas to about 260° C. The product phosphoric acid (68-76% P₂O₅) is splitfrom this recycle stream. The remaining phosphoric acid is removed fromthe gases leaving the spray tower in a high pressure drop venturi, whichhas recirculated, cooled weak phosphoric acid as the scrubbing liquidused as required to make up liquid volume in the spray tower liquid. Theproduct scrubber in FIG. 1 includes the spray tower and venturi and mayproduce any concentration of P₂O₅ less than or equal to about 76%.However, 72% may be selected because it is at the bottom of a eutecticand remains more fluid at room temperature, albeit still with arelatively high viscosity.

An induced draft (ID) fan pulls the gasses through the kiln and off gasscrubbing system of FIG. 1. The gasses exiting the ID fan are treatedwith a flue gas desulfurization scrubber to remove acidic componentsfrom the off gas before it discharges to the environment.

IHP Example 1

Raw phosphate ore from the Meade formation near Soda Springs, Id. wascombined with raw silica ore from a mine near Soda Springs, Id. anddelayed green petroleum coke from a gulf coast petroleum refinery suchthat the calcium-to-silica mole ratio was 0.56 and the fixed carbon inthe petroleum coke was 1.3 times the theoretical to chemically reducethe phosphate in the mixture. The mixture contained 13.43 wt % P₂O₅,21.33 wt % CaO, 42.42 wt % SiO₂, and 7.38 wt % fixed carbon on a drybasis. The mixture was dried and then ground to 80.2%-200 mesh in anopen circuit ball mill. Twelve percent water was added to the groundmixture and formed in a die into several ½ inch diameter pellets whichwere then dried in an oven at 280° F. One pellet was placed in agraphite crucible with a graphite lid and the crucible and its contentsheated at 20° C./min to 1180° C. and held for one hour under anatmosphere of nitrogen. After cooling the crucible to room temperaturethe pellet was analyzed to determine how much phosphorus had evolvedfrom the pellet. The experiment was repeated with the time at 1180° C.increased to two hours and then to four hours. It was found that 61%,82% and 97% of the phosphorus had evolved from the pellet in 1, 2, and 4hours at 1180° C., respectively.

Over 100 experiments were additionally performed using the sameprocedure with a range of temperatures, impurity levels, reaction times,calcium-to-silica mole ratios, grind sizes, and carbon addition levels.Some of the most significant variables are summarized in FIG. 2.

IHP Example 2

Several of the spent pellets from IHP Example 1 were analyzed andcompared to the analyses of the feed pellets to determine the fate ofimpurities in the feed compositions. The elements that volatilizedeither partially or totally from the pellet during the phosphatereduction together with the weight fraction of the element in the feedpellet, were Ag (<1 ppm), As (10.2 ppm), C (8.2%), Cd (33 ppm), Cl (90ppm), Cs (1.3 ppm), F (1.27%), Ga (3 ppm), Hg (0.22 ppm), K (0.44%), N(0.13%), Na (0.25%), P (6.55%), Pb (7 ppm), Rb (18 ppm), S (0.73%), Se(6.8 ppm), Tl (1 ppm), and Zn (688 ppm). The elements which were presentin significant amounts, which were of concern, and which evolved fromthe feed pellets were K, Na, S, and F. FIG. 3 shows the weight fractionof these elements which volatilized from the pellet over a range of holdtemperatures as compared to the weight of the elements in the feedagglomerates. The composition of the pellets tested in thisvolatilization study is the same as given in the 1180° C. tests given inIHP Example 1.

Thus, in a fully reducing environment, part of these elementsvolatilized from the kiln solids. To the extent an oxidizing shell formsfrom carbon burnout, the loss of these elements is expected to be muchhigher therein.

IHP Example 3

The co-ground ore/silica/green petroleum coke mixture indicated in IHPExample 1 above was blended with 9 wt % moisture and formed into +¼/−½inch balls on a 3 foot balling pan with an additional 3 wt % moistureadded through control sprays on the balling pan. The balls were dried ina tray drier, and a twenty pound charge of balls was processed in a 13.5inch diameter by 22¼ inch long batch rotary kiln at 1320° C. The kilnwas fired with 1 to 2 feet³/minute (cfm) of propane, about 44 cfm ofair, and 0 to 2 cfm of oxygen as necessary to hold the temperature atthe set point. The kiln was fitted with an onboard thermocouple and hadzircon refractories.

The kiln was preheated to 1189° C. and twenty pounds of balls werecharged to the batch kiln with a scoop. Samples of the kiln solids weresampled every fifteen minutes and analyzed to determine the burnout ofcarbon and the extent of phosphate reduction. The surface of the kilnbed reached 1271° C. by the minute sample, overshot to 1332° C. by the30 minute sample, and was 1320° C. for the 45 minute sample. The carboncontent of the entering ball was 8.75%, but the sum of volatiles andcarbon that reacted with iron oxide and carbonate during heat up wasestimated at 0.96% leaving 1.56 pounds of fixed carbon in the 20 poundcharge to the furnace. The amount of phosphorus that had evolved, thecarbon which had burned out, and the carbon remaining in the balls isgiven for the three samples in Table 2.

TABLE 2 Carbon burnout and phosphorus evolved from 13.5 inch batch kilntest. Sample P evolved C Reacted C Burnout C Remaining 15 minute 10%  8%49% 43% 30 minute 27% 21% 62% 17% 45 minute 29% 23% 75%  2%

From this data 62% of the fixed carbon that was in the ball was burnedout by oxidizing gasses in 16 minutes as the bed solids heated from 700°C. to 1180° C. where the phosphate reduction reaction commenced. Thesurface area of the bed was 1.69 square feet and the volume of the bedwas 0.28 cubic feet. Therefore the carbon burnout rate was 3.46 poundsper hour per square foot, and the heat up time times thesurface-to-volume ratio equals 97 min-ft²/ft³.

The twenty pounds of feed agglomerates charged to the kiln containedgreen petroleum coke with 0.19 pounds of volatiles. Since thesevolatiles emerged from the bed over about a fifteen minute period, theirflux was about 0.45 pounds per hour per ft². The phosphorus and carbonmonoxide gases from the run were 29 wt % of the total phosphate ore masscharged to the batch kiln and totaled 0.67 pounds over about twentyminutes or about 1.19 pounds per hour per ft². The carbon burnout andphosphate reduction were occurring so rapidly with regard to the 15minute samples in the very small batch kiln that the protective effectof the volatiles from the green coke was not sufficient to be clearlyobserved. Nevertheless, the existence of a protective effect of thephosphorus metal and carbon monoxide was suggested in the data from thistest.

Samples of the product balls after 45 minutes in the kiln were split intwo and a white shell and black core with a sharp boundary between themwas observed. The thickness of the shell was measured at a number ofpoints around the circumference on several different balls to obtain anaverage thickness. From these measurements and the diameter of the ball,the volume fraction of the shell with the carbon burned out wasdetermined to be 70%, and quite consistent over several balls. The shellwas determined to contain silica, Ca₃(PO₄)₂, and an amorphous phaselikely to be calcium metaphosphate. The phosphorus in the shell wasanalyzed at 22.5 wt %, which is considerably enriched over the startingphosphorus level in the feed ball. The core contained less than 1 wt %residual phosphorus. The shell composition indicated the reaction ofphosphorus metal in the shell or the area between the balls withoxidizing gases from the freeboard, forming P₄O₁₀ which, in turn,reacted with fluorapatite in the shell to form the amorphous phase.Table 3 lists analyses of the feed ball, shell, and core.

TABLE 3 Analysis of the core and shell from the 45 minute sample inbatch kiln test of IHP Example 3. Component Feed Ball Core Shell P₂O₅13.95% 0.3%* 20.7% CaO 20.16% 23.25%  18.42% MgO 0.279% 0.324% 0.240%SiO₂ 24.6% 42.33%  35.34% K₂O 0.326% 0.390% 0.149% F 1.54% 0.647% 0.057%S 0.783% 0.488% 0.022% *From XRD analysis.

IHP Example 4

A mixture of washed, but not beneficiated phosphate ore matrix directfrom a Florida phosphate mine and calcined petroleum coke with ananalysis of 11.0% P₂O₅, 15.9% CaO, 56.2% SiO₂, and 11.4% fixed carbonrespectively, was ground to 80% −325 in a dry ball mill and formed intoballs using a balling pan. These balls had a calcium-to-silica moleratio in the feed of 0.304 and fixed carbon at 2.45 times thetheoretical amount to chemically reduce the phosphate ore therein. Thedried balls were passed into a 33 inch diameter (inside the bricks) by30 foot long counter-current kiln at 600 pounds per hour, ultimatelyreaching a bed temperature of 1520° C. near the burner end of the kilnat the furthest point into the kiln that could be observed with anoptical pyrometer. The dense, white flame of burning phosphorus obscuredmaking temperature measurement further into the kiln.

During the continuous kiln operation, the kiln was stopped briefly andsamples of kiln solids were taken at the discharge of the kiln and witha long probe at 3¾ foot, 7½ feet, 13¼ feet, and 15 feet from the hot endof the kiln. The samples were analyzed for carbon, calcium, andphosphorus. The percent of phosphorus and carbon volatilized, the mass(lbs/hr) of carbon consumed in the phosphorus reduction (C(P rx)), andthe mass (lbs/hr) of carbon burned out (CBO) at each sample location wascalculated with the results shown in Table 4.

TABLE 4 Analytical data from continuous KPA pilot plant run. lbs C Distfrom Feed % P loss % C loss (P rx)/hr* lbs CBO/hr Feed 0 0 0 0 15 feet 220.2 0.6 13.3 18.75 feet 10 35.7 2.8 21.6 22.5 feet 49 52.6 13.7 22.326.25 feet 65 58.6 18.1 22.0 30 feet 72 61.4 20.1 21.9 This is theamount of carbon in pounds/hr that was consumed by reacting tochemically reduce phosphate ore within the ball.

The kiln was not ported and all of the oxygen required to oxidize thephosphorus and carbon monoxide arising from the bed was burned with airadded at the burner end of the kiln. The observed position wherephosphorus was volatilized given in Table 4 showed that most of thephosphorus and carbon monoxide were generated and burned over a narrowlength of the kiln at the burner end of the kiln. Thus, a narrowtemperature spike occurred with the peak temperature of 1520° C. to getthe 72% phosphorus yield during the short time duration of thetemperature spike. The kiln overall had an off gas that was reducing,meaning that it contained no excess oxygen. Often, reducing off gas thuscontains unoxidized CO and may even contain some unoxidized P₄.

The kiln solids had a relatively long heat up time where the kiln solidswere between 700° C. and 1180° C., during which time carbon was burnedout of the shell. The calcined petroleum coke used in the feedformulation did not evolve any volatiles to protect the kiln bed duringheat up. However, once significant generation of the phosphorus metaland carbon monoxide evolved from the kiln solids commenced about 19 feetinto the 30 foot kiln, the rate of carbon burnout was essentiallystopped as shown in the lbs CBO/hr column of Table 4.

The amount of phosphorus and carbon monoxide that evolved per hour inthe kiln was 40.8 pounds per hour, which evolved over about a 6 footlength of the kiln in which the solids spent about 30 minutes and whichhad a bed surface area of 13.37 ft². In this case the resulting flux of6.1 pounds per hour per square foot substantially limited oxidizing gastransport to the bed and essentially stopped carbon burnout during theperiod of phosphate reduction. The product calcium metasilicate glassyresidue also encapsulated the remaining carbon and helped protect itfrom carbon burnout.

The spent balls that exited the kiln were broken in half and observed tohave a white shell and a black interior with a sharp line of demarcationbetween the two areas. The white shell had no fixed carbon and highlevels of phosphorus remaining. The black core had phosphorus reductionof over 96% and carbon remaining in the core.

The heat up time of the bed from feed temperature to 1180° C. was about2.23 hours, which gives a bed surface-to-volume ratio times heat up timevalue of 334 minutes-ft²/ft³. The carbon burnout rate during heat up was0.272 pounds per hour per square foot.

Hypothetical IHP Example 5

The observations made in the preceding IHP examples allow calculation ofthe performance of a full-scale commercial KPA kiln run according to thevarious embodiments of the IHP. Kiln feed balls made from raw matrix orewith calcium-to-silica mole ratio of 0.5 and a fixed carbon content of1.3 times the theoretical amount to chemically reduce the phosphate inthe ore are made by combining 1 part raw matrix phosphate ore with 15%contained P₂O₅ from the Mead formation in Idaho, trammeled to remove +3inch oversize material, and 0.09 parts green petroleum coke sized 3 inchand less. This mixture is dried in a rotary drier, sized to ⅜ inch andless and co-ground in an open circuit ball mill to 80%-200 mesh. Theground ore is formed into +¼/−½ inch balls on a balling drum which aredried in a grate drier. One and a half percent coal sized to +¼/−½inches is added to the balls prior to the grate drier.

This feed is processed at the rate of 490,000 pounds per hour in a 19foot diameter (inside the bricks) by 220 foot long rotary kiln equippedwith a porting system to supply over bed air as required to provide thedesired temperature profile, on board brick thermocouples, and an offgas analyzer capable of analyzing CO, CO₂, and O₂. The loading in thekiln is 15% and the kiln burden spends one and one/half hour in thekiln. It is heated to 1260° C. in 30 minutes and the temperature held at1260° C. until the solids discharge from the kiln one hour later. Thebed volume is 9,945 ft³ and the bed surface area is 3,600 ft².

The coal which has 25% volatiles which produces about 880 pounds perhour of volatiles while the incoming kiln feed mixture is heated from anentering temperature of 125° C. to 600° C. which occurs in 10 minutes.Thus, the flux of hydrocarbons in the cold end of the kiln is 2.3 poundsper hour per ft², which by comparison with the phosphate reduction fluxthat provided protective above, is sufficient to give a modestprotective effect to prevent P₄O₁₀ from the freeboard from condensing onthe cold balls.

Heating the balls from 600° C. to 950° C. takes another 10 minutes.During this time the volatiles from the green petroleum coke and thecoal total 4,480 pounds providing a protective flux of 45 pounds perhour per square foot. In addition, the light molecular weight of thevolatiles and their ability to react with the incoming oxidizing gasesmake them more effective than phosphorus metal and carbon monoxidefluxes which were found to be effectively protective at a flux one ninthas great in IHP Example 4.

During heating from 950° C. to 1180° C. which takes an additional 10minutes the balls are subject to carbon burnout at a rate of about 1pound per hour per square foot over 1660 square feet or about 1660pounds per hour of carbon burnout. The sacrificial carbon interceptsabout 40% of the incoming oxidation so the fixed carbon burned out ofthe balls is about 1000 pounds per hour or about 1/40th of the fixedcarbon in the incoming balls. This results in the carbon being burnedout of a thin shell on the outside of the feed balls, which is notchemically reduced after the balls reach reaction temperature. Thus aloss in yield of just over 2% results from carbon burnout in this area.

The bed volume to surface area ratio times the heat up time in thiscommercial kiln is 11 minutes-ft²/ft³.

After the balls reach 1180° C. until they discharge from the end of thekiln a total of 98% of the phosphate ore is chemically reduced givingrise to 91,800 pounds per hour of phosphorus metal and carbon monoxideflux through 6,630 square feet of bed or a flux of 13.8 pounds per hourper square feet ore several times the flux found in IHP Example 4 to besubstantially protective against carbon burnout. During this time thefixed carbon from the coal on the outside of the balls and the carbonrich core inside the ball allow some reduction of phosphorus in theshell, mitigating somewhat the loss of yield from carbon burnout in theshell. Three percent of the phosphorus that forms remains in the bed inthe form of ferrophos due to the relatively high amount of iron found inmatrix ore.

Since the carbon burnout is now at a low level the reduction of thephosphate in the balls can be allowed to take place at an optimumtemperature which is sufficient to react the phosphate during the onehour dwell time in the hot zone of the kiln, but low enough to reducethe loss of sodium, potassium, fluorine, and sulfur to the freeboardwhile the phosphorus is removed from the kiln residue. Operation at thisrelatively low temperature, but over an extended length of the kiln,also gives high throughput, long refractory life, freedom from problemsassociated with melting in the kiln burden, and freedom from problemsassociated with sinoite formation in the kiln. The relatively lowoverall carbon burnout allows operation of the kiln with an oxidizingoff gas so that an afterburner is avoided.

Thus, this IHP example provides a process wherein relatively cheap greenpetroleum coke provides the reactant and heat to produce high qualityphosphoric acid product (68-76% P₂O₅) at the rate of 523,000 pounds P₂O₅per hour from as mined matrix ore and to produce a kiln residue that isinsoluble and can be used for landfill.

According to a further embodiment of the IHP, a phosphorous pentoxideproducing method includes forming a kiln bed using feed agglomeratesand, initially, from about 1 to about 2 wt % bedding coke or coal in acounter-current rotary kiln, the bed having a length and having asurface-to-volume ratio of less than about 1.15 ft²/ft³ and the kilnhaving an inner diameter of at least 6 feet. The agglomerates containphosphate ore particles, carbonaceous material particles containing fromabout 8 to about 12 wt % volatiles, and sufficient silica particles forthe agglomerates to exhibit a calcium-to-silica mole ratio of from 0.5to less than 1.0, at least 80% of all particles exhibiting a size lessthan 200 mesh. Individual agglomerates substantially exhibit a sameelemental composition, a same calcium-to-silica mole ratio, and a sameproportion of excess fixed carbon at least 1.3 times a theoreticalcarbon requirement for reduction of all phosphate in the ore.

The method includes maintaining the bed surface-to-volume ratio with aninlet dam sufficiently tapered such that it reduces turbulence comparedto otherwise identical processing with an untapered dam. The method alsoincludes adding over bed air and/or oxygen through a plurality of portsalong the bed length, the ports being of sufficient location, number,and throughput to decrease a time for bed heat up to 1180° C. comparedto otherwise identical processing without the over bed air and/oroxygen. The method further includes maintaining a bed temperature at orabove 1180° C. along at least 50% of the bed length without exceeding1310° C. along the entire bed length. The method still further includesgenerating kiln off gas and collecting phosphorous pentoxide from thekiln off gas, the kiln discharging a residue containing processedagglomerates, less than 10% of the agglomerates' phosphate input to thekiln remaining in the residue as phosphate. By way of example, less than5% of the input phosphate might remain in the residue. Calculation ofthe maximum surface-to-volume ratio for this embodiment of the IHPassumes a kiln length of 72 ft and a solids loading of 15%. Generally,the aspect ratio of the kiln may be about 12 or less (i.e., 6 ftdiameter×12=72 ft). The various embodiments of the IHP herein may allowfor a ratio as low as about 10.

In compliance with the statute, the invention has been described inlanguage more or less specific as to structural and methodical features.It is to be understood, however, that the invention is not limited tothe specific features shown and described, since the means hereindisclosed comprise preferred forms of putting the invention into effect.The invention is, therefore, claimed in any of its forms ormodifications within the proper scope of the appended claimsappropriately interpreted in accordance with the doctrine ofequivalents.

The invention claimed is:
 1. A phosphorous pentoxide producing methodcomprising: forming a reducing kiln bed using feed agglomerates below anoxidizing freeboard of a counter-current rotary kiln, the bed having alength; the agglomerates containing phosphate ore particles,carbonaceous material particles, sufficient silica particles for theagglomerates to exhibit a calcium plus magnesium-to-silica mole ratio ofless than 1.0, and a lignin sulfonate both inside the agglomerates andcoating a surface of individual agglomerates; maintaining a bedtemperature at or above 1180° C. along a portion of the bed length;generating freeboard particulates from the agglomerates, an amount ofparticulates generated being less than would occur in the method with nolignin sulfonate; and generating kiln off gas and collecting phosphorouspentoxide from the kiln off gas, the kiln discharging a residuecontaining processed agglomerates, less than 20% of the agglomerates'phosphate input to the kiln remaining in the residue, and the percentageof input phosphate that remains in the residue being less than wouldoccur in the method with no lignin sulfonate.
 2. The method of claim 1wherein less than 10% of the agglomerates' phosphate input to the kilnremains in the residue and the calcium plus magnesium-to-silica moleratio is less than 0.5.
 3. The method of claim 1 further comprisingforming green feed agglomerates from a homogeneous blend of water, thelignin sulfonate, the phosphate ore particles, the carbonaceous materialparticles, and the silica particles, the green agglomerates containingabout 0.1 to 4 wt % lignin sulfonate.
 4. The method of claim 3 furthercomprising controlling a drying rate of the green agglomerates and, asthe water evaporates from the agglomerates, forming the lignin sulfonatecoating on individual dried agglomerates by wicking some of the ligninsulfonate onto the surface of individual drying agglomerates withoutadhering the agglomerates together.
 5. The method of claim 1 wherein thecarbonaceous material particles comprise about 8 to 12 wt % volatilesand at least 80% of all particles exhibit a size less than 200 mesh. 6.The method of claim 1 wherein the lignin sulfonate coating has athickness of about 1 to 50 micrometers and individual agglomeratescontain a total of about 0.1 to 4 wt % lignin sulfonate including thecoating.
 7. The method of claim 1 wherein the lignin sulfonate insidethe agglomerates exhibits a continuously decreasing concentrationgradient of lignin sulfonate from the surface into individualagglomerates, individual agglomerates comprising a total of about 0.1 to4 wt % lignin sulfonate including the coating.
 8. The method of claim 1further comprising volatilizing at least a portion of the ligninsulfonate coating during bed heat up from about 650 to about 1180° C. 9.The method of claim 1 further comprising blanketing the bed withvolatiles evolving at least partly from the lignin sulfonate during bedheat up from about 120 to about 650° C. and retaining a sacrificiallayer of fixed carbon from the lignin sulfonate coating.
 10. The methodof claim 9 further comprising volatilizing at least a portion of thesacrificial layer during bed heat up from about 650 to about 1180° C.11. A phosphorous pentoxide producing method comprising: co-grinding amixture containing phosphate ore, carbonaceous material, and silica;homogeneously blending a lignin sulfonate with the co-ground mixture;forming green feed agglomerates from the blended lignin sulfonate andco-ground mixture; drying the green agglomerates and wicking some of thelignin sulfonate onto a surface of individual drying agglomerates, whichforms a lignin sulfonate coating thereon, the dried agglomeratescontaining phosphate ore particles, carbonaceous material particlescontaining about 8 to 12 wt % volatiles, and sufficient silica particlesfor the dried agglomerates to exhibit a calcium plus magnesium-to-silicamole ratio of less than 0.5 and at least 80% of all particles exhibitinga size less than 200 mesh; forming a reducing kiln bed using the driedagglomerates below an oxidizing freeboard of a counter-current rotarykiln, the bed having a length; blanketing the bed with volatilesevolving at least partly from the lignin sulfonate during bed heat upfrom about 120 to about 650° C. and retaining a sacrificial layer offixed carbon from the lignin sulfonate coating; volatilizing at least aportion of the sacrificial layer during bed heat up from about 650 toabout 1180° C.; maintaining a bed temperature at or above 1180° C. alonga portion of the bed length; and generating kiln off gas and collectingphosphorous pentoxide from the kiln off gas, the kiln discharging aresidue containing processed agglomerates, less than 10% of theagglomerates' phosphate input to the kiln remaining in the residue. 12.The method of claim 11 wherein the homogeneous blend further includeswater and the drying comprises controlling a drying rate of the greenagglomerates and, as the water evaporates from the agglomerates, formingthe lignin sulfonate coating on individual dried agglomerates by wickingsome of the lignin sulfonate onto the surface of individual dryingagglomerates without adhering the agglomerates together.
 13. The methodof claim 11 wherein the green feed agglomerates comprise green feedballs exhibiting a greater than ¼ to less than ½ inch diameter prior todrying and the dried feed balls exhibit a dry crush strength greaterthan 25 pounds.
 14. The method of claim 11 wherein the lignin sulfonatecoating has a thickness of about 10 to 40 micrometers and a remainder ofthe lignin sulfonate exhibits a continuously decreasing concentrationgradient of lignin sulfonate from the surface into individualagglomerates, individual agglomerates comprising a total of about 0.4 to2 wt % lignin sulfonate, including the coating.
 15. The method of claim11 further comprising generating freeboard particulates from thephosphate ore particles, an amount of particulates generated being lessthan would occur in the method with no lignin sulfonate.
 16. The methodof claim 11 wherein the percentage of input phosphate that remains inthe residue is less than would occur in the method with no ligninsulfonate.
 17. Phosphate ore feed balls produced by the processcomprising: co-grinding a mixture containing phosphate ore, carbonaceousmaterial, and silica; homogeneously blending water and a liquidcontaining a lignin sulfonate with the co-ground mixture; forming greenfeed balls from the blended water, lignin sulfonate, and co-groundmixture, the green balls containing about 0.1 to 4 wt % ligninsulfonate; drying the green balls by passing heated air through amultilayer bed of green balls; and controlling a drying rate of theballs and, as water evaporates from the balls, forming a ligninsulfonate coating on individual dried balls by wicking some of thelignin sulfonate onto a surface of individual drying balls withoutadhering the balls together, a remainder of the lignin sulfonate inindividual dried balls forming a continuously decreasing concentrationgradient of lignin sulfonate from the surface into individual driedballs, the dried balls containing phosphate ore particles, carbonaceousmaterial particles containing about 8 to 12 wt % volatiles, andsufficient silica particles for the dried balls to exhibit a calciumplus magnesium-to-silica mole ratio of less than 1.0 and at least 80% ofall particles exhibiting a size less than 200 mesh.
 18. The phosphateore feed balls of claim 17 wherein the green feed balls exhibit agreater than ¼ to less than ½ inch diameter prior to drying and thedried feed balls exhibit a dry crush strength greater than 25 pounds.19. The phosphate ore feed balls of claim 17 wherein the carbonaceousmaterial comprises green petroleum coke and the calcium plusmagnesium-to-silica mole ratio is less than 0.5.
 20. A phosphate orefeed ball comprising: a dry, homogeneous mixture of phosphate oreparticles, carbonaceous material particles containing about 8 to 12 wt %volatiles, and sufficient silica particles for the ball to exhibit acalcium plus magnesium-to-silica mole ratio of less than 1.0, at least80% of all particles exhibiting a size less than 200 mesh; a ligninsulfonate both inside the ball and coating a surface of the ball, thecoating having a thickness of about 1 to 50 micrometers and a remainderof the lignin sulfonate exhibiting a continuously decreasingconcentration gradient of lignin sulfonate from the surface into theball, the ball containing a total of about 0.1 to 4 wt % ligninsulfonate, including the coating, and exhibiting a dry crush strengthgreater than 25 pounds.
 21. The phosphate ore feed ball of claim 20wherein the carbonaceous material comprises green petroleum coke and thecalcium plus magnesium-to-silica mole ratio is less than 0.5.